Search Documents

By Timothy Batchler

"Standing roof supports help to provide a stable working environment in coal mine gateroads. NIOSH participates in the development of these support systems by conducting full-scale tests in their Mine Roof Simulator. These tests have established the performance characteristics of every coal mine support system currently in use in the United States. Through these tests, mine operators and regulatory officials have acquired a fundamental understanding of the various types of support performance that guides the assessment and use of these support systems. However, some little known factors that influence support behavior can lead to misconceptions and, perhaps, a poor assessment of support behavior at times. This paper highlights several of these factors and discusses their significance. The paper also examines a current longwall tailgate support and discusses the inefficiency with this approach. Finally, the NIOSH Mine Roof Simulator (MRS) is the world’s most powerful load frame built specifically to test full-scale roof supports, including longwall shields. are addressed in this paper to provide a more in-depth assessment of the limitations of the MRS and what the MRS testing protocol is designed to do. INTRODUCTION Over the years, considerable research has been conducted to develop standing support technologies and to evaluate their performance characteristics. The load and displacement characteristics of the standing supports have been determined from full-scale testing conducted using the National Institute for Occupational Safety and Health (NIOSH) Mine Roof Simulator (MRS) (Barczak and Tadolini, 2008) located in Bruceton, PA. Performance traits, installation patterns, and ground responses are observed in various geological and mining conditions and are evaluated to determine the support performance in the field (Dolinar, 2010; Campoli, 2015; Zhang et al., 2012). Despite widespread use of standing supports, not every aspect of their performance and resulting ground support capability are always well understood. This paper discusses 10 factors, several of which are surprising, that impact standing support performance."

Jan 1, 2017

By Brian E. Perrine

Underground mining requires a variety of specialized equipment. With ever-changing mining conditions and methods, some mines have had to resort to purchasing existing industrial equipment and adapting it for underground use. Since this equipment is not made for underground mining or their specific task, the result is often lost time, money, and sometimes an unsafe working condition. One such application is the setting of a CAN. A CAN is a type of cylindrical secondary roof support made from a cement composite and encased in steel. The CAN has become popular because of its simple one-piece design that can handle heaving and sagging better than other available designs. ?Because of this exceptional ability to sustain load through extended convergence and displacement, the CAN is one of the most robust and stable standing supports in use today.?It also promotes better air flow due to the size and shape of the CAN. CAN size and weight have increased over time to handle more difficult mining conditions. Significant down time and inefficiencies with existing CAN setting machines at Twentymile Coal were directly due to the item in use not being designed to be task-specific. They were undersized with very limited movement. The miners had to take extra steps and precautions when setting CANs. Maintenance delays were significant. Twentymile Coal Company realized that they required a safer, more reliable and more efficient machine to handle and set CANs, and looked for a partner to develop a new design-specific machine. They worked jointly with J.H. Fletcher® engineering and sales to design and manufacture a machine specifically made for setting CANs. The objective was to produce a machine that would handle the conditions and methods that the mine required for setting CANs during longwall retreat. The resulting design had to be safe, stable, have a wide range of motion, be ergonomic, simple to use, and easy to maintain. The machine also had to improve efficiency. The resulting machine has been operating for approximately one year at Twentymile Coal and has been proven to improve safety, reliability, and efficiency. In conclusion, with the task-specific CAN Manipulator, the mine is able to better plan daily tasks with the knowledge that the machine will be available throughout a shift, and the employees will be safer because of the machine?s ergonomics, stability and safety. This has resulted in better productivity and a safer environment.

Jan 1, 2013

By Guy Reed, Russell Frith, Kent Mctyer

"Coal pillar design has historically been based on assigning a design factor of safety (FoS) or stability factor (SF) to coal pillars according to their estimated strength and the assumed overburden load acting on them. Acceptable FoS VALUES (have been assigned based on past mining experience, and at least one methodology includes the determination of a statistical link between FoS and probability of failure (PoF).The role of the pillar width-to-height (w/h) ratio has long been established as having a material influence on both the strength of a coal pillar and its potential mode of failure. However, there has been significant professional disagreement on using both FoS and w/h ratio as part of a combined pillar system stability criterion, as compared to using FoS in isolation. The argument is that, as w/h ratio is intrinsic to pillar strength, which, in tum, is intrinsic to FoS, it makes no sense to include w/h ratio twice in the stability assessment. At face value, this logic is sound.However, this paper will argue that there is a valid technical reason to bring w/h ratio into system stability criteria (other than its influence on pillar strength), as it is related to the post-failure stiffness of the pillar, as measured in situ, and its interaction with overburden stiffness.When overburden stiffness is bought into pillar system stability considerations, two issues emerge. The first is the width-to-depth (W/H) ratio of the panel, in particular whether it is sub-critical or super-critical from a surface subsidence perspective. As a minimum, this directly relates to the accuracy of the pillar loading assumption of full tributary area loading. The second relates to a reevaluation of pillar FoS based on whether the pillar is in an elastic or non-elastic (i.e., post-yield) state in its as-designed condition, as relevant to maintaining overburden stiffness at the highest possible level.The significance of the model presented is the potential to maximise both reserve recovery and mining efficiencies without any discernible increase in geotechnical risk, particularly in thick seams and higher depth of cover mining situations. At a time when mining economics are, at best, marginal, removing potentially unnecessary design conservatism without negatively affecting safety is of interest to all mine operators and is an important topic for discussion amongst the geotechnical community."

Jan 1, 2016

By Kourosh Shahriar, Kazem Oraee, Ezzeddin Bakhtavar

"In this paper, a relationship between dependant parameters and the crown pillar thickness is first introduced. This relationship defines geotechnical problems caused by thin pillars and economic considerations created by pillars that are thicker than the optimum size. For this purpose, a dimensional analysis as an effective physico-mathematical tool was used. This technique restructures the original dimensional variables of a problem into a set of dimensionless products using the constraints imposed upon them by their dimensions. A model is hence introduced that calculates the optimum pillar thickness. The relationship introduced here and the method applied can be used by mining engineers in all situations where a combined open-pit and block caving method is deemed to be the most appropriate mining method.INTRODUCTIONMany deposits can be mined entirely with the open-pit method; others must be worked underground from the very beginning. In addition, there are the near-surface deposits with considerable vertical extent. Although they are initially exploited by the open pit method, there is often a ""transition depth"" where decision has to be made about changing to underground methods.Some of the biggest open-pit mines worldwide will reach their final pit limits in the next 10 to 15 years. Furthermore, there are many mines planning to change from open-pit to underground mining due to increasing extraction depths and environmental requirements. In this way, it is likely that block and/or panel caving will enable the operations to continue achieving a high production rate at low costs as an underground method (Bakhtavar et al., 2009).In these cases, it is often necessary to consider a crown pillar beneath the transition depth (open-pit floor) before starting an underground caving stope method (Figure 1).There are generally two kinds of crown pillar: a ""surface crown pillar"" and ""crown pillar between open-pit and underground mining."" In general, they both play the similar role in mining. Since the primary purpose of a surface crown pillar is to protect surface land users, the mine, and those working in it from inflows of water, soil, and rock; it is vital that the surface crown pillars remain stable throughout their life. When a crown pillar could potentially remain between open-pit and underground mining it is specially proposed to prevent water entering from the open-pit floor into the stope, as well as to reduce open-pit wall and floor caving."

Jan 1, 2010

By Ross Seedsman

There are many models for how roof bolts behave, one of which is prevention of shear along bedding. The basis of this model is that a roof beam will not delaminate if bedding parallel shear can be prevented. A key point in such a model is that bedding discontinuities are present naturally in the roof and that mining induced stresses do not generate new surfaces. The presence of these discontinuities means that the rock mass must be modelled as a transversely isotropic material if continuum codes are used. A limit-equilibrium type design method for specifying roof bolts is being developed using voussoir beam and excess shear force concepts. The method incorporates much of what operators know intrinsically about roof reinforcement, and allows mining engineers to focus on the key uncertainty - the frequency and location of bedding surfaces in the immediate roof. Areas where more research is required are highlighted.

Jan 1, 2000

By Michael J. Castner

"INTRODUCTIONMultiple-seam mining is an ever growing concern for engineers and mine management in every portion of the mining industry and is especially true for coal mining operations. The goal of this report is to identify the location and severity of multiple-seam interactions, provide AMSS calculations performed on the worst of these locations, and follow up with a visual examination of the underground physical effects on the roof and ribs in the locations identified for analysis. Multiple-seam mining is increasingly being studied because:1. Multiple-seam mining can create a multitude of problems such as roof falls, pillar crushing, pillar failing, rib spalling, floor heave, and inundation of water, gas, or oxygen-deficient air that can disrupt mining operations and threaten the safety of the miners.2. Multiple-seam mining is becoming more prevalent due to the depletion of virgin coal reserves.Currently, coal mining in the United States has progressed to the point that a growing number of mining companies are operating in coal seams located above, below, and/or adjacent to previously mined out coal seams. As a result, the state and federal governments are enacting higher safety standards on mining operations. For example, in years past, pillars may have been properly designed to handle single seam interactions for a particular mine or mines; however, today, where mining is now being considered above or below a previously mined area, these loading conditions may not be adequate to handle the removal of additional seams above or below the previously mined reserve. Studies from 1981 estimated that two-thirds of the mineable coal reserves in the United States would be impacted by multiple-seam implications. Today, over three decades later, it is obvious that most of the remaining mineable coal reserve will be influenced by previously extracted seams. Thus, thefocus on multiple-seam mining interactions on current and future mining operations. In 2012, underground mining in the United States produced 342,287 thousand short tons of coal. Northern Appalachia accounted for over 30% (103,301 thousand short tons) of this total (U.S. Energy Information Administration, 2013). Rosebud Mining Company and Western Allegheny Energy (RMC and WAE) have a combined total of 23 actively producing underground mining operations and at least four new mines under or ready for development. All of these mines are located in coal seams (Upper and Lower Freeport; Upper, Middle, and Lower Kittanning) of northern Appalachia (Pennsylvania and Ohio). Currently, more than half of the RMC and WAE mines are dealing with multiple-seam interactions. Seven mines are undermining previously mined out coal seams, and five mines are overmining previously mined out coal seams. Also, between these companies, five mines have inner-seam sloping possibilities (2 completed, 3 planned), most of which will result in mining over or under previously mined areas."

Jan 1, 2015

By Christopher Mark

"Many problems in rock mechanics are limited by our imperfect knowledge of the material properties and failure mechanics of rock masses. Mining problems are somewhat unique, however, in that plenty of real world experience is generally available and can be turned into valuable experimental data. Every pillar that is developed, or stope that is mined, represents a full-scale test of a rock mechanics design. By harvesting these data, and then using the appropriate statistical techniques to interpret them, mining engineers have developed powerful design techniques that are widely used around the world. Successful empirical methods are readily accepted because they are simple, transparent, practical, and firmly tethered to reality.The author has been intimately associated with empirical design for his entire career. But where his past papers have described the application of individual techniques to specific problems, the focus of this paper is the process used to develop a successful empirical method. A six-stage process is described:1. Identification of the problem, and of the end users of the final product2. Development of a conceptual rock mechanics model, and identification of the key parameters in that model3. Identification of measures for each of the key parameters, and the development of new measures (such as rating scales) where necessary4. Data sources and data collection5. Statistical analysis6. Packaging of the final productEach of these stages has its own potential rewards and pitfalls, which will be illustrated by incidents from the author’s own experience. The ultimate goal of this paper is to provide a new and deeper appreciation for empirical techniques, as well as some guidelines and opportunities for future developers."

Jan 1, 2015

By Chase K. Barnard, Rahul Thareja, Sean Warren, Rajagopala Kallu

"Inflatable rock bolts are commonly utilized for ground support in Nevada underground mines, however, there is limited information regarding what factors influence the bond strength of these bolts. Bond strength is an important parameter in friction bolt support design, and this information is usually not available from manufacturers as a standard. Previous research has investigated the bond strength of friction bolts; however, these studies focused primarily on split set bolts and ground conditions more favorable compared to those commonly encountered in Nevada underground mines. Without site-specific bolt pull-out tests, ground control engineers are required to make assumptions regarding the in-situ bond strength of rock bolts. This is especially a concern for projects in the feasibility and initial development stages before site-specific experience is acquired.The purpose of this paper is to establish confidence in anticipated minimum bond strength for inflatable rock bolts by comparing the bond strength to variable geotechnical conditions using the Rock Mass Rating (RMR) system. To investigate a correlation between these parameters, the minimum bond strength of pull-out tested inflatable rock bolts was compared to the RMR of the rock in which these bolts were placed. Bond strength vs. RMR plots indicate that expected minimum bond strength is positively correlated with RMR, however the correlation is not strong. Cumulative percent graphs indicate that 97% of pull-out tests result in a minimum bond strength of 1 ton/ft and ½ ton/ft in RMR =45 and <45, respectively.Although lower bond strengths are more commonly encountered in low RMR ground, high bond strengths are possible as well, yielding higher variability in bond strengths in low RMR ground. Bond strength of friction bolts relies on contact between the rock bolt and drill hole. Experience in Nevada indicates that RMR is known to affect both the quality and consistency of drill holes which likely affects bond strength. Drilling and bolting in low RMR ground is more sensitive to drilling and bolting practices, and strategies for maximizing bond strength in these conditions are discussed."

Jan 1, 2015

By Kevin Andrews, Scott Nelson, Terry Hamrick

"In the process of Geologic Hazard Mapping (GHM) for underground coal mining, geologists evaluate the effects of various geologic and geotechnical conditions to assist with optimization of mine productivity and with selection of more favorable areas for future development. The methodology involves correlation of multiple geologic factors (such as overburden depth, proximity to sandstone channels, and areas with weak immediate roof strata) with conditions encountered during mining (roof falls, floor heave, pillar spalling, etc.). Such studies yield maps that identify areas where mining is likely to be difficult and allow mining engineers to foresee favorable and unfavorable mining areas when planning new or expanding mines. To expand the benefits of GHM, work has been done to correlate geologic factors with room-and-pillar mining advance rates. Average mining advance rates are mapped and then correlated to different areas, as characterized by GHM. This process exposes the geological characteristics that have a material impact on mining rates of advance, improving the rates of advance into future mining areas with similar geological conditions. The final result of the work allows mine planners to not only design the mine qualitatively accounting for areas of unfavorable conditions, but also allows the mine planner to more accurately apply rates of advance for mine timing and production sequencing. The use of GHM to predict mining advance rates in unmined areas enhances the benefits of traditional hazard mapping and provides quantitative data that can be incorporated into mine sequencing models.INTRODUCTIONThe role of a geologist in the coal mining industry has evolved significantly over the last 30 to 40 years. Geologic Hazard Mapping (GHM) mapping for coal mines began in the late 1970s and early 1980s (Chase, Newman, and Rusnak, 2006). Since its beginnings, the GHM process has evolved significantly and continues to take on new levels of sophistication. Numerous case studies of GHM application to predict future mining conditions exist in the literature. For example, Artrip, Nelson, and Newman (1993), documents the geotechnical characterization of roof and floor for mining in the Imboden Seam in Virginia where previous mining efforts had been abandoned due to adverse mining conditions. In addition, Keim and Miller (1999) summarize evaluation of geological influences on a longwall mine in West Virginia. In both examples, significant geotechnical and geological information was available, including the opportunity to compare predicted mining conditions against actual conditions as mining progressed."

Jan 1, 2018

By Long Fan, Shimin Liu, Guijie Sang

"Weak shale roof falls have long been the main cause for fatalities in underground coal mines. Moisture-induced swelling is one of the root causes for the degradation of roof shales. In this study, slake durability tests are conducted on shale samples collected from Illinois basin coal mines, and different levels of durability are analyzed according to Gamble slake durability classification. There are two main mechanisms influencing the shape of retained fragments after two cycles of wetting and drying: mechanical abrasion and swelling pressure. A prototype optically based shale swelling apparatus is designed and manufactured for the shale free-swelling measurement. The experimental results suggest that swelling pressure tends to deteriorate shale laminations. Moisture-induced swelling of the No. 6 roof shale from Bear Run Mine was measured under 100% relative humidity condition. The measured swelling strain normal to beddings is ~5 to 7 times greater than the swelling strain parallel to beddings. This suggests that interlayer expansion plays a key role in moisture-induced swelling. Cracks induced by swelling pressure tend to occur along the bedding plane and lead to shale deterioration. INTRODUCTION Background Ground falls have long been the cause for nearly 50% of all fatalities in underground coal mines and continue to be one of the greatest safety hazards in underground spaces. Among the various ground fall hazards, roof failure takes up 23% of total fatalities in underground coal mines, of which 13% are due to skin failure and 10% are due to massive roof falls (Mark, Pappas, and Barczak, 2011). Roof failures associated with weak shales are summarized in previous studies (Molinda and Mark, 2010; Aughenbaugh and Bruzewski, 1973; Bajpayee, Pappas, and Ellenberger, 2014; Murphy, 2016). For mine roof instability and failure, two aspects can explain the failure mechanism of weak roofs. One is the elevated stress concentration in lateral direction due to the mining-induced excavation and its stress redistribution around the mine opening (Song and Stankus, 1997; Fan and Liu, 2017; Hill, 1986). The other mechanism is the passive stress-relief due to the strength weakening induced by the ambient moisture and temperature fluctuations during a period of time. Shale is moisture sensitive and presents a lower durability if it is exposed to a highly humid environment for a long time, thus leading to roof deterioration and instability in underground coal mines (Aughenbaugh and Bruzewski, 1976; Stateham and Radcliffe, 1978; Chugh and Missavage, 1981; Klemetti, Oyler, and Molinda, 2009). It is well known that shale is a clay-bearing sedimentary rock and that clay minerals are water sensitive. The moisture uptake for shale will induce the micro-deformation of a shale matrix, thus the structure and strength of shale is a function of time and intensity of moisture exposure. All mine engineers notice moisture-induced shale deterioration; however, the underlying mechanism(s) are still not fully understood due to the complexity of geophysical-chemical water-shale interactions (Huang, Aughenbaugh, and Rockaway, 1986; Hensen and Smit, 2002; Diaz-Perez, Cortés-Monroy, and Roegiers, 2007). The coal mining industry is still in the process of finding a technically sound, cost-effective technology to prevent this unexpected and unpredictable shale roof failure. Therefore, understanding how the moisture-sensitive shale responses to various ambient moisture and what role exposure duration plays in shale roof failure is key to prevent unexpected roof incidents, which will lay the foundation for the long-term roof support design and ground control management."

Jan 1, 2017

By Pinnaduwa H. S. W. Kulatilake, Biao Shu

The intact rock properties and discontinuity properties for both Devonian Rodeo Creek (DRC) and Devonian Papovich (DP) rock formations that exist in the selected open-pit mine were determined from tests conducted on rock samples. Special survey equipment, which has a total station, laser scanner and a camera, was used to perform remote fracture mapping in the research area. From remote fracture mapping data, the fracture orientation, spacing and density were calculated using more refined procedures compared to what exist in the literature. Discontinuity orientation distributions obtained through remote fracture mapping agreed very well with the results of manual fracture mapping conducted by the mining company. The GSI rock quality system and Hoek-Brown failure criteria were used to estimate the rock mass properties combining the fracture mapping results with laboratory test results of intact rock samples. Fault properties and the DRC-DP contact properties were estimated based on the laboratory discontinuity test results. A geological model was built in a 3DEC model including all the major faults, DRC-DP contact and two stages of rock excavation. The built major discontinuity system of 44 faults in 3DEC with their real orientations, locations and three-dimensional extensions were validated successfully using the fault geometry data provided by the mining company using seven cross-sections. Numerical modeling was conducted to study the effect of boundary conditions and lateral stress ratio on the stability of the considered rock slope. For the considered section of the rock slope, the displacements obtained through stress boundary conditions seemed more realistic than those obtained through zero velocity boundary conditions (on all four lateral faces). Stable deformation distributions were obtained for k0 in the range of 0.4 to 0.7. Because the studied rock mass is quite stable, it seems that an appropriate range for k0 for this rock mass is between 0.4 and 0.7. The displacements occurred between July 2011 and July 2012 due to the nearby rock mass excavation that took place during the same period; they were compared between the field monitoring results available from the mining company and the predicted numerical modeling results and a good agreement was obtained. Overall, the successful simulation of the rock excavation during a certain time period indicated the possibility of using the procedure developed in this study to investigate rock slope stability with respect to expected future rock excavations in mine planning.

Jan 1, 2015

By Bruce Hebblewhite, Jim Galvin

"A coal burst occurred on 15 April, 2014 at the Austar Coal Mine, located west of Newcastle, NSW, Australia. The burst resulted in fatal injuries to two men working as part of the mining crew at the development face. At the time, a continuous miner was being used to mine a longwall development gate road through heavily structured coal, at a depth of approximately 550m. A number of pre-cursor bumps had occurred on previous shifts, emanating from the coal ribs of the roadway, in proximity to the coal face.This paper reviews the geological, geotechnical and mining conditions and circumstances leading up to the coal burst event; and presents and discusses the available evidence and possible interpretations relating to the geomechanical behaviour mechanisms that may have been critical factors in this incident. The paper also discusses some key technical and operational considerations of ground support systems and mining practices and strategies needed for operating in such conditions in the future.INTRODUCTIONAustar Coal Mine is an underground longwall coal mine located near Cessnock in the Hunter Valley of New South Wales (NSW), Australia and is the only underground mine still extracting the Greta Seam in this region. The mine was the first in Australia to adopt the Chinese-developed Longwall Top Coal Caving (LTCC) method for thick seam extraction. Typically seam thickness ranges from 4m to 7m and depth of mining from 480m to 560m, with future mining planned down to depths of up to 700m, making it one of the deepest operating coal mines in Australia.On 15 April 2014, a pressure burst occurred in the left hand rib at the active mining face of B Heading, 2 to 3 cut-through, Maingate A9 panel, during development of the gateroads for the ninth longwall top coal caving panel. Strata in the general vicinity was affected by disturbed geology and multiple geological structures. Figure 1 shows a section of the mine plan as at October 2014 (six months after the accident), indicating the current longwall extraction panel (AS) and the development panel A9 where the accident occurred. (Note: Neither development nor longwall face positions changed significantly between the time of the accident and the date of this plan). At the time of the accident the current longwall face was in excess of 1,000m away from the Maingate A9 development panel face position."

Jan 1, 2016

By Syd Peng, Gamal Rashed

"Bumps in coal mines have been recognized as a major hazard for many years. These sudden violent failures around mine openings have compromised safety, ventilation and access to mine workings. Previous studies showed that the violence of coal specimen failure depends on both the interface friction and width-to-height (W/H) ratio of a coal specimen. The mode of failure for a uniaxially loaded coal specimen or a coal pillar is a combination of both shear failure along the interface and compressive failure in the coal. Unstable interface failures provide a sudden de-confinement to coal rib, thus triggering uns and compressive failure in the sidewalls. The compressive failure of a coal specimen or a coal pillar can be controlled by changing its W/H ratio. As the W/H ratio increases, the ultimate strength increases. With a proper combination of interface friction and the width-to-height (W/H) ratio of pillar or coal specimen, the mode of failure will change from sudden violent failure, which is brittle failure to non-violent failure, which is ductile failure. The main objective of this paper is to determine at what W/H ratio and interface friction the mode of failure changes from violent to non-violent. In this research, coal specimens of width-to-height ratio ranging from 1 to 10 were uniaxially tested under two interface frictions of 0.10 and 0.25. The results are presented and discussed.INTRODUCTIONCoal mine bumps are sudden, violent bursts of coal from a pillar or pillars or a block of coal, resulting in a section, the whole pillar, or the solid block of coal being thrown into an open entry. These bursts are accompanied by very loud noises (Peng, 2008). Several case histories in mining have long described coal pillars or coal faces failing violently with an accompanying ejection of debris and broken material into the working areas of the mines (Campoli et al., 1987; Osterwald, 1962; Peperakis, 1958; Garvey et al., 2013). Because of the catastrophic nature of these sudden failures, understanding the causes of coal mine bumps is essential to create a safe underground working environment. This phenomenon has motivated many ground control researchers to conduct extensive field investigations throughout the past century. Rashed and Peng (2014) compared the mechanical properties of two kinds of coal, one from bump and the other from non-bump-prone mines. They found that coal itself does not play any significant role in coal bumps and recommended that future research should focus more on the local variation of geological conditions such as the interface friction between pillar and roof and between pillar and floor."

Jan 1, 2015

By Marvin B. Thompson

"The depositional model for major Illinois Basin coalbeds is well documented and defined on a regional scale. The model was a driving force during past exploration for lower sulfur and thicker coal deposits. The most attractive coal reserves were known to lie along the flanks of contemporaneous sandstone filled stream channels and were overlain by thick silty gray shale roof. At various distances away from the channels, the depositional model transitioned to black shale I limestone roof, sulfur content of the coal bed increased, and the coal bed often thinned. Understanding the character of roof rock types associated with this depositional model is essential to predicting ground control factors on both a regional scale, and on resolving roof control problems at the mine-site. Additionally, horizontal stress and overburden depth are significant factors that impact roof rock quality and ground control in Illinois Basin mines.Determining the character of several rock types, and defining the primary horizontal stress direction were crucial to a recent successful project at New Future Mine, located in southern Illinois. An inclined slope was constructed from the Illinois No. 5 Coalbed up to the Illinois No. 6 Coalbed through strata that ranged from extremely weak ""stackrock"" and a weak underclay unit, to a massive sandstone unit. Engineering geology, operations management, and Jennmar technical services personnel teamedup to select and install roof support for each strata transition. The project was completed with no injuries and no roof falls.INTRODUCTIONAbundant published information concerning coal deposition, geographic distribution, quantity, quality, and mining conditions exists for Illinois Basin coal reserves. The State of Illinois, through its State Geological Survey (ISGS), has provided decades of research, record keeping, and service to the coal industry. Other excellent sources of Illinois Basin coal information include the Indiana Geologiq1l Survey and the Kentucky Geological Survey. Past research by the U.S. Bureau of Mines (USBM) supplemented the work by the geological surveys, especially in the area of horizontal stress and mining hazards. The National Institute for Occupational Safety and Health (NIOSH) continues to produce publications focused on roof control issues common to many coal basins, including the Illinois Basin. Many of the illustrations in this paper originated in ISGS and USBM publications."

Jan 1, 2010

By B. Macdonald

In 1991, roofbolts were introduced as a primary support for inseam retreat longwall entries at CBDC&apos;s Phalen Colliery. Increasing mine depth (+550 m) necessitated replacing conventional inseam steel sets in these entries because they were being severely damaged during wall retreat and subsequently causing gate end problems. Roofbolts had been used in the Sydney coalfield between 1949 and 1963 in room and pillar working, however, some concerns arose with using roofbolts as a primary support in long (2.2 - 3.4 km) single entry roadways. In order to alleviate these concerns, a phased approach was adopted whereby conventional steel densities remained along with roofbolts. The roofbolt design was such that the roofbolts would be the primary support and the steel sets would experience little or no loading. A detailed geotechnical monitoring program was developed by both CBDC and the Cape Breton Coal Research Laboratories technical staffs. The results of this monitoring were used to build-up confidence in roofbolts and subsequently reduce the steel set densities until eventually only roofbolts were used as a primary support. This paper will discuss briefly the phased introduction of rootbolts at Phalen Colliery and more specifically the development of geotechnical data collection and interpretation.

Jan 1, 1994

By S. P. Singh

Destress blasting is commonly used in the mines to minimize the occurrence of rockbursts. The fracturing created by destressing in the vicinity of the face reduces the stress concentration by pushing the zone of high stress deeper into the rock mass. But the destressing practice by blasting may be devastating near the faults or other sensitive high stress zones. To avoid severe rockbursts in such situations, the application of chemical destressing has been proposed in this paper. A laboratory investigation was conducted on the rocks in the Sudbury basin to examine the change in the material properties with an aim of assessing the feasibility of destressing by chemical means. The favourable change observed in the rock properties led to the suggestion that the injection of surfactants in boreholes would enhance fracture propagation similar to what is currently done by blasting but in a safer, inexpensive and non-violent manner. The current destressing practices in Ontario mines have also been reviewed in this paper.

Jan 1, 1987

By Tulu B. Ihsan, Mark A. Van Dyke, Ted M. Klemetti

"How well do current modeling procedures calculate the rear abutment extent and loading? Does an improved understanding of the rear abutment extent warrant a change in standing support in bleeder entries? What is the optimal standing support for bleeder entries separated from the longwall startup room by a barrier pillar? To help answer these questions and to determine the current utilization of standing support in bleeder entries, four bleeder entries at varying distances from the startup room were instrumented, observed, and numerically modeled. This evaluation was intended to determine the rear abutment extent and magnitude at various locations to optimize standing support in these entries and those under similar conditions. This paper details observations made by NIOSH researchers in the bleeder entries of a deep cover longwall panel; specifically data collected from instrumented pumpable cribs, observations of the conditions of the entries, and numerical modeling of the bleeder entries during longwall extraction. The primary focus was on the extent and magnitude of the abutment loading experienced by the standing support. As expected, the instrumentation of the standing supports showed very little loading relative to the capacity of the standing supports—less than 25 tons load and 1 inch convergence. The observations of the conditions showed little to no change from before the longwall panel extraction began to when the panel was more than 50% extracted. In addition to the observation and instrumentation, numerical modeling was performed to evaluate the bleeder design. The Flac3D program was used to evaluate these four bleeder entries using previously defined modeling and input parameter estimation procedures. The results indicated only a minor increase in load during the extraction of the longwall panel. The model showed a much greater increase in stress due to the development of the gateroad and bleeder entries, with about 80% of the increase associated with development and 20% with longwall extraction. The Flac3D model showed very good correlation between expected gateroad loading during panel extraction and those expected based on previous studies. The front and side abutment extent modeled was very similar to observations from this and previous panels with similar conditions. The results of this study showed that the rear abutment stress experienced by this bleeder entry design was minimal. The convergence measured in these bleeder entries was small enough not to mandate any additional support beyond development even by U.S. or Australian standards. The pumpable crib load and convergence experienced depends on the geological setting, mining geometries, and timing. The farther away from the startup room, the lower the applied load and smaller the convergence in the entry if all else is held constant. Finally, the numerical modeling method used in this study was capable of replicating the expected and measured results near seam."

Jan 1, 2017

By Michael Jenkins

This paper reviews the history of rock burst research in the Coeur d&apos;Alene Mining District of northern Idaho over the last 30 years. Rock bursting is a problem because of a combination of geological factors in the region--complex structure, high horizontal stresses, and strong, brittle rock that readily stores strain energy. The Bureau of Mines has cooperated with industry to study rock burst phenomenon for many years. Early work focused on methods to mitigate bursting problems within the context of existing mining methods. In the 1970&apos;s, researchers began investigating methods for altering or controlling conditions that lead to rock bursts. These control measures become part of the stope development that precedes mining or are changes in the mining process itself. Although, researchers continue to develop microseismic techniques, the major research effort focuses on developing a new mining method to reduce the number and severity of rock bursts in order to improve safety and increase productivity in the Coeur d&apos;Alene Mining District.

Jan 1, 1987

By Heather E. Lawson

Recent field data providing measured pillar loading from a deep western mine revealed that an increased load transfer distance existed compared to expected values (Larson et al., 2012). Additionally, recent work by Tulu and Heasley (2012) has shown that the abutment angle in deep longwall mines can be significantly reduced as the depth of mining increases. The load transfer distance and abutment angle are both used as parameters to estimate pillar loading in longwall design. To investigate the potential implications of these observations on pillar loading in deep longwall mines, a parametric study using ALPS loading estimation procedures was performed. A supplemental study was conducted using LaModel to further investigate pillar loading at the tailgate. The study found that the abutment angle is particularly important for panels with supercritical geometries. As panels reach a subcritical geometry with increasing depth, the primary influence of the abutment angle is reduced and matched by other factors, including load transfer distance for tailgate loading. An extended load transfer distance was shown to potentially reduce loading on pillars adjacent to the longwall as the loading is spread over a larger area. The results of these studies indicate a need for further field verification and research toward regional geologic influences on pillar loading and how these may impact ground stability.

Jan 1, 2013

By Sandin E. Phillipson

"On April 5, 2010 a massive dust-fueled explosion at a longwall mine in southern West Virginia claimed the lives of 29 miners. The mine had experienced large gas inrushes from the floor on two known previous occasions, and natural gas was found emanating from floor fractures behind the longwall shields after the explosion, indicating a likely fuel source for the initial ignition. In succeeding months, another longwall mine operating in the same seam 15 miles away encountered smaller inflows of natural gas that halted production from two different longwall faces. Production was halted at a third longwall mine, 30 miles south of the first, in May 2011, when explosive levels of methane were detected inby the longwall face on the headgate. The same mine was evacuated on March 20, 2014 when a series of floor gas feeders were encountered on the longwall face. A literature review indicated that there has been little documentation or description of these events, and that their size and frequency are unknown. The most significant floor gas inrushes in the United States are believed to have been associated with longwall mining in the Pocahontas No. 3 Seam of western Virginia, in which an estimated 5.8 x 107 ft3 of natural gas was expended from a floor feeder on the longwall face over a two-week period (Aul, pers. comm.).A variety of mechanisms and circumstances are associated with methane emanations from underground mines. Hyman (1987) defined an outburst as a violent, simultaneous release of gas and comminuted rock material into a working face or the interior of a borehole, and indicated that certain soft or fractured coals have a propensity for outburst based on gas desorption rates. In this mechanism, reservoirs of autogenic gas hosted in pockets of soft, crushed coal surrounded by harder coal are released suddenly as confining stress is removed by mining. Hyman (1987) also suggested that highly stressed, relatively gassy coal that is penetrated by mining-induced fractures could be outburst-prone. Clayton et al. (1993) and Ulery (2008) recognized the importance of faults as potential conduits for gas from adjacent strata into coal mines, and suggested that biogenic gas trapped in adjacent strata could migrate into mined coal seams due to a pressure differential. Ulery (2008) distinguished between an outburst and a blower, the latter defined as an event expending a large amount of gas over an extended time period, most often emanating from underlying strata, but without expulsion of coal or rock. Methane feeders are further defined as a subset of blowers that continually expend gas over a long period of time but at a lower rate. Many of the mechanisms discussed by Ulery (2008) relate to geologic features that can impound gas within the seam, releasing a voluminous amount when mining breaches the elastic dike, fault gouge, or other geologic structure that had restricted and compartmentalized coalbed methane migration. Thus, the events discussed in this paper can be described as blowers and feeders. However, detailed geological observations of the three longwall mines discussed in this paper appear to indicate a mechanism that is different from those described by Hyman (1987) or Ulery (2008), in that they did not involve coalbed-sourced gas, and the reservoirs were not hosted by coal seams."

Jan 1, 2016